Process for the recovery of value metals from base metal sulfide ores

ABSTRACT

A process for leaching a value metal from a base metal sulfide ore, comprising the step of leaching the ore with a lixiviant comprising a chloride, an oxidant and hydrochloric acid is disclosed. The leaching is controlled, by use of low concentrations of hydrochloric acid and a redox potential, to effect formation of hydrogen sulfide from the base metal sulfide ore. The hydrogen sulfide is stripped from the leach solution, thereby reducing the amount of sulfate generated in the leach to very low levels. The leaching may also be conducted to limit the co-dissolution of platinum group metals and gold with the base value metals. The leach forms a value metal-rich leachate and a solids residue. The solids residue may be subsequently leached to recover the platinum group metals and gold. The value metal-rich leachate can be is oxidized and neutralized to recover the value base metals. In an embodiment, the chloride is magnesium chloride and lixiviant solution is regenerated.

FIELD OF THE INVENTION

The present invention relates to a method for the leaching and recoveryof value metals, especially nickel, copper, zinc and cobalt values, andPlatinum Group Metals (PGMs) and gold from base metal sulfide ores,including from mixtures of sulfide and oxide ores. In particularembodiments, the base metal sulfide ores are value metal-containing oresor concentrates, especially pyrrhotite, pentlandite, chalcopyrite,arsenopyrite and other pyrites, sphalerite, and concentrates and mattesthereof. The leaching may be conducted using a low concentration ofhydrochloric acid, in chloride media. In particular, the method may beoperated such that sulfide in the ore is substantially converted tohydrogen sulfide, and preferably essentially converted to hydrogensulfide, rather than to sulfate or to elemental sulfur. In preferredembodiments, the hydrogen sulfide formed is stripped from the leachsolution, thereby providing leachate with a low concentration of sulfurand/or sulfate. Conversion of the sulfide of the ore to, in particular,hydrogen sulfide simplifies and/or allows for alternate steps forseparation and recovery of value metals.

BACKGROUND OF THE INVENTION

Base metal sulfide ores exist in many areas of the world and are apotential source of many value metals. In particular, the ores maycontain zinc, nickel, copper, cobalt and the PGMs, silver and gold. Theprincipal ores are all iron-bearing, and examples particularly includenickeliferous pyrrhotite Fe₈S₉, pentlandite (FeNi)₉S₈, chalcopyriteCuFeS₂, arsenopyrite FeAsS and sphalerite ZnS. Cobalt may be found inthe lattice of a pentlandite ore. Base metal sulfide ores have been usedextensively in the commercial production of nickel, cobalt, zinc andcopper.

Base metal sulfide ores may be processed using hydrometallurgical orpyrometallurgical techniques. Recovery of nickel, copper and PGMs tendsto be high with the pyrometallurgical route, typically being greaterthan 90%, and cobalt recovery is typically between 30 and 70%. Recoveryof nickel, cobalt, zinc and copper is also high in thehydrometallurgical route, but PGMs and gold tend to be lost in the leachresidue unless further, often complicated and costly, recovery processesare carried out.

Smelting of nickel sulfide concentrates produces a liquid furnace matte.The liquid furnace matte is then subjected to air oxidation, in aprocess known as converting, to remove most of the iron and sulfur. Ironand gangue impurities are removed as a disposable slag. The resultingconverter matte, also known simply as matte, may then be treated toobtain the nickel, cobalt, copper and PGMs and gold. The treatmentmethods used are mainly hydrometallurgical, for example refiningprocesses based on sulfate, carbonyl, ammoniacal and chloride chemistry.Sulfate and especially chloride-based refining processes are discussedby G. Van Weert in “Some Observations on the Chloride Based Treatment ofNickel-Copper-Cobalt Mattes” pages 277-298 of Chloride Metallurgy2002—Volume 1, 32^(nd) Annual Hydrometallurgy Meeting, Edited by E. Peekand G. Van Weert, published by CIM.

In a chloride leach process, the most valuable component, viz. nickel,may be solubilized first, with little leaching of copper, thus achievinga separation of nickel from copper. In a known chloride leach process(Thornhill, P. G., Wigstol, E and Van Weert, G., “The Falconbridge MatteLeach Process”, Journal of Metals, 23(7), 1971 p13) using very stronghydrochloric acid, the leach may be represented as follows:Ni₃S₂+6HCl=>3NiCl₂+2H₂S+H₂

In an alternative leach process based on chlorine, a granulatedconverter matte is ground and fed to a chlorine leach process where itis subjected in a first step to a redox controlled leach processsolubilizing most of the nickel and part of the copper, but none of thePGMs:Ni₃S₂+3Cl₂=>3NiCl₂+2S⁰Cu₂S+Cl₂=>CuS+CUCl₂

To remove cupric copper, which is the regarded as the leachant,additional matte is added without chlorine, followed by cementation. Inanother alternative, advantages of a chlorine leach could be achievedusing sub-azeotropic hydrochloric acid and oxygen. Solubilized copper(cupric) chloride would again be the leaching agent.

The hydrometallurgy of complex sulfide bulk concentrates is discussed byD. S. Flett in “Chloride Hydrometallurgy for Complex Sulfides: a Review”pages 255-276 of Chloride Metallurgy 2002—Volume 1 above. In particular,the ferric or cupric chloride leaching of Cu/Pb/Zn/Ag type sulfideconcentrates is discussed. Recent activity in the treatment of singlesulfide concentrates, particularly copper e.g. pressure leaching usingBrCl₂ ⁻ as oxidant, is also reported. The article concludes that this isthe most promising process for commercialization but the development ofprocessing of complex sulfide concentrates still has some way to gobefore commercialization is finally realized.

A process for recovering non-ferrous metal values from ametal-containing sulfide material containing at least one of zinc,copper, lead, cobalt, nickel, silver and gold, as well as iron, isdisclosed in U.S. Pat. No. 4,378,275 of Adamson et al, issued Mar. 29,1983. The sulfide material is leached under oxidizing conditions withacidic aqueous chloride lixiviant solution containing magnesiumchloride. The oxidizing conditions are disclosed as use of molecularoxygen in the form of air, oxygen-enriched air and pure oxygen. Althoughleaching at atmospheric pressure is stated to be possible, it ispreferable to operate the leach stage under elevated partial pressures,i.e. under pressure leach conditions. Use of elevated temperatures ispreferred, i.e. at least about 50° C. to about 250° C., withtemperatures in the range of 100° C. to 180° C. being preferred. Theperiod for leaching is from about 5 minutes to about 12 hours. The useof low chloride levels is preferred. For example, Adamson et al.provides that the chloride ion concentration is typically from about 4to about 6 grams of ions per liter. The kinetics of the process wouldindicate a need to use long periods of leaching at the lowertemperatures and atmospheric pressure. Pressure leaching, using oxygen,of a Zn/Cu/Fe ore containing very low levels of nickel at 160° C. isexemplified. In the process, non-ferrous metal values are solubilized,leaving iron oxide and sulfur as a residue. The leach liquor issubjected to liquid—liquid extraction using a hydrophobic extractant.The raffinate, containing magnesium chloride and any sulfates formedduring the leach process, is subjected to pyrohydrolysis to yieldhydrogen chloride and magnesium oxide. The sulfates are then removed bywashing of the magnesium oxide formed, which counteracts many of theadvantages of forming magnesium oxide by pyrohydrolysis.

SUMMARY OF THE INVENTION

In one aspect, this invention provides a method for the separation ofsulfur, which is derived from sulfides associated with base metals, froma lixiviant produced during the leaching of base metals from a basemetal sulfide ore or concentrate, especially nickel, copper, zinc,cobalt and PGMs, silver and gold, that operates at atmospheric pressure.In accordance with this aspect, the process relates to a method for thereduction of the amount of sulfur in the leachate and leach solids byseparation of sulfide in the ore as hydrogen sulfide during the leachingof a base value metal from a base metal sulfide ore. In particular,sulfur may be removed from the leachate and the solid leach residue byforming and stripping hydrogen sulfide in the leaching step. Some or allof the hydrogen sulfide may subsequently be converted to elementalsulfide. An advantage of this process is that sulfur in the form ofhydrogen sulfide may be separated by simple gas/liquid separationtechniques. The hydrogen sulfide may be used for downstream purificationtreatments of the leachate and/or as a relatively pure source ofhydrogen sulfide for use in the production of sulfur compounds, such aselemental sulfur.

In another aspect, this invention provides a process for the recovery ofvalue metals from base metal sulfide ore, concentrate or matte byleaching with a lixiviant having a high chloride concentration and a lowconcentration of hydrochloric acid. The use of a lixiviant with a highchloride loading permits the use of lower concentrations of hydrochloricacid in the lixiviant.

In has now been determined that by adjusting the redox potential and thepH of a lixiviant, base metals associated with a sulfide may be leachedfrom a sulfide source material using a lixiviant having a high chloridecontent and a relatively low concentration of hydrochloric acid whilePGMs and gold are essentially not leached and wherein a substantialportion, and preferably essentially all, of the sulfide that is leachedis converted to hydrogen sulfide. It will be appreciated that, in someembodiments, it may be determined to leach some of the PGMs and goldwith the base metals. In other embodiments, it may be determined toconduct the leach of the base metals so that a portion of the sulfidesulfur that is dissolved is not converted to hydrogen sulfide. Theextent to which some of the PGMs and gold may be leached with the basemetals and the sulfide sulfur that is dissolved and converted tohydrogen sulfide will vary depending upon several factors including thecomposition of the sulfide source material, the degree of metal recoverythat is selected and the reaction kinetics that are selected for theleaching step.

In accordance with one aspect of the present invention, there isprovided a process for leaching a value metal from a sulfide orematerial containing said value metal, said sulfide ore material,comprising the step of leaching the sulfide ore material at atmosphericpressure with a lixiviant comprising hydrochloric, a chloride selectedfrom the group consisting of alkali metal chlorides, magnesium chlorideand calcium chloride, and mixtures thereof, and an oxidant selected fromthe group consisting of alkali metal peroxide, alkali metal perchlorate,ammonium perchlorate, magnesium perchlorate, alkali metal chlorate,alkaline earth metal perchlorate, chlorine, alkali metal hypochlorite,hydrogen peroxide and peroxysulfuric acid, and mixtures thereof toobtain a leachate and a solid residue.

In one embodiment, the process further comprises selecting a sulfide orematerial that contains at least one value metal selected from the groupconsisting of nickel, copper, zinc and cobalt, and mixtures thereof.

In another embodiment, the process further comprises selecting a sulfideore material that additionally contains at least one of gold and aplatinum group metal.

In another embodiment, the process further comprises selecting a sulfideore material that comprises a base metal sulfide ore, a concentrate of abase metal sulfide ore, a matte obtained from a base metal sulfide ore,tailings from the processing of a base metal sulfide ore and mixturesthereof.

In another embodiment, the process further comprises selectingpyrrhotite, pentlandite, chalcopyrite, pyrite, arsenopyrite andsphalerite, and mixtures thereof as the base metal sulfide.

In another embodiment, the process further comprises selecting sodiumchloride, potassium chloride, magnesium chloride and calcium chloride,and mixtures as the chloride.

In another embodiment, the process further comprises selecting magnesiumchloride as the chloride.

In another embodiment, the process further comprises selecting chlorine,sodium chlorate, hydrogen peroxide, sodium hypochlorite and sodiumperchlorate and mixtures thereof as the oxidant.

In another embodiment, the process further comprises selecting sodiumchlorate, chlorine and mixtures thereof as the oxidant.

In another embodiment, the process further comprises adjusting the pHand the redox potential during the leach so that the pH is less than 2.5and the redox potential is in the range of 200 to 600 mV.

In another embodiment, the process further comprises adjusting the redoxpotential during the leach so that the redox potential of the leachsolution is in the range 250-450 mV.

In another embodiment, the process further comprises selecting theconcentration of magnesium chloride to be at least 200 g/L, preferablyto be in the range of 200-500 g/L and more preferably to be in the rangeof 200-400 g/L, the total concentration being formed essentially frommagnesium chloride and hydrochloric acid. Optionally, the amount ofhydrochloric acid is in the range of 30-150 g/L.

In another embodiment, the process further comprises conducting theleach at a temperature in the range of from 75° C. to the boiling pointof the solution at ambient pressure.

In another embodiment, the process further comprises adjusting the pH sothat, at the end of the leach, the pH is less than 1.5 and, preferablyless than 1.

In another embodiment, the leachate and a solid residue are subjected toa solids/liquid separation step and the solid residue is subjected to afurther leaching step to recover at least a portion of the gold and atleast one of the platinum group metal.

In another embodiment, the solid residue is also treated for separationof a magnetic fraction.

In another embodiment, the process further comprises selecting the oreor concentrate to comprise values of nickel and iron, and one or more ofcobalt and copper.

In another embodiment, the process further comprises:

-   -   (a) subjecting the leachate to a series of value metal recovery        steps and obtaining a value metal depleted leachate and    -   (b) treating the value metal depleted leachate to recycle at        least a portion of the hydrochloric acid and chloride.

Step (b) preferably comprises pyrohydrolysis. Preferably, step (a)includes at least one precipitation step using a base. Preferably thebase is magnesium oxide. Preferably, magnesium oxide is produced fromthe value metal depleted leachate.

In another embodiment, the process further comprises treating theleachate to precipitate iron.

In another embodiment, the leachate is treated by increasing the pH ofthe leachate, subsequent to removal of the residual solids to, toprecipitate iron.

In accordance with another aspect of the present invention, there isprovided a process for leaching a value metal from a sulfide sourcematerial containing said value metal with a lixiviant comprisinghydrochloric, a metal chloride, and an oxidant, the leaching with saidlixiviant being controlled so that at least about 50% of sulfide sulfurthat is leached from the sulfide source material is converted tohydrogen sulfide during said leaching to obtain a leachate and a solidresidue.

In one embodiment, the leaching is controlled so that at least about 90%of sulfide that is leached from the sulfide source material is convertedto hydrogen sulfide during said leaching.

In another embodiment, the leaching is controlled so that at least about99% of sulfide sulfur that is leached from the sulfide source materialis converted to hydrogen sulfide during said leaching.

In another embodiment, the process further comprises selecting thesulfide source material to contain at least one value metal selectedfrom the group consisting of nickel, copper, zinc and cobalt, andmixtures thereof and to additionally contain at least one of gold andplatinum group metal and the leaching is controlled so that the gold andplatinum group metals are essentially not leached.

In another embodiment, the process further comprises selecting thesulfide source material to contain at least one value metal selectedfrom the group consisting of nickel, copper, zinc and cobalt, andmixtures thereof and to additionally contain at least one of gold andplatinum group metals and the leaching is controlled so that less than10 weight percent of the gold and platinum group metals are leached.

In another embodiment, the leaching is controlled by adjustment of thepH and the redox potential.

In another embodiment, the process further comprises adjusting the redoxpotential of the lixiviant to be from 250 to 600 mV.

In another embodiment, the process further comprises adjusting the pH ofthe lixiviant at the end of the leach to be less than 2.5.

In another embodiment, the process further comprises adjusting thelixiviant so as to have a concentration of chloride ions of from 200-500g/L.

In another embodiment, the process further comprises selecting theamount of hydrochloric acid that is added to obtain a selected pH.

In another embodiment, the process further comprises selecting thesulfide source material from a base metal sulfide ore or a materialderived from a base metal sulfide ore.

In another embodiment, the chloride comprises an alkali metal chloride,magnesium chloride and calcium chloride, and mixtures thereof, and theoxidant comprises an alkali metal peroxide, alkali metal perchlorate,ammonium perchlorate, magnesium perchlorate, alkali metal chlorate,alkaline earth metal perchlorate, chlorine, alkali metal hypochlorite,hydrogen peroxide and peroxysulfuric acid, and mixtures thereof.

In another embodiment, the process further comprises conducting theleach at atmospheric pressure.

In accordance with another aspect of the present invention, there isprovided a process for leaching a value metal from a sulfide orecomprising leaching the sulfide ore with a lixiviant comprisinghydrochloric acid, a metal chloride, and an oxidant to obtain a leachateand a solid residue, wherein the redox potential is maintainedsufficiently low to essentially not leach platinum group metals and goldfrom the sulfide ore material and to essentially convert the sulfidesulfur which is leached from the sulfide ore to hydrogen sulfide.

In one embodiment, the sulfide ore comprises a base metal sulfide ore, aconcentrate of a base metal sulfide ore, tailings from the processing ofa base metal sulfide ore and mixtures thereof. Preferably, the sulfideore is unroasted prior to being leached. Preferably, the leach isconducted at atmospheric pressure.

In another embodiment, at least 90% by weight of the sulfide sulfur thatis leached from the sulfide ore material is converted to hydrogensulfide.

In another embodiment, the process further comprises the step oftreating at least some of the hydrogen sulfide to obtain elementalsulfur.

In another embodiment, the pH of the lixiviant at the end of the leachis less than 2.5, preferably less than 1.5, more preferably less than 1and most preferably less than 0.

BRIEF DESCRIPTION OF THE DRAWINGS

The present invention will be described with reference to the preferredembodiments of the invention shown in the drawings, in which:

FIG. 1 shows a flow sheet for the recovery of value metals fromsulfide-based nickeliferous ore or concentrate; and,

FIG. 2 shows an alternate flow sheet for the recovery of value metalsfrom sulfide-based nickeliferous ore or concentrate.

DETAILED DESCRIPTION OF THE INVENTION

The present invention relates to a process for the leaching of a valuemetal from a base metal sulfide source material. The base metal sulfidesource material may be present with a base metal oxide source material.For example, the source material may be a mixture of sulfide andoxide-based ores.

The ores may be an ore per se, but is preferably a concentrate thereof.In other embodiments, the ore may be in the form of any of the mattesdiscussed above, especially converter matte, or in the form of tailingsof a base metal sulfide ore. It is understood that the expression “ore”also includes an ore and any materials derived from an ore. Preferably,the ore is unroasted.

In embodiments of the invention, the ore comprises, and preferablyconsists essentially of, an ore known as pyrrhotite, pentlandite,chalcopyrite, arsenopyrite, sphalerite, a pyrite and mixtures thereof.As noted above, the ore may be a mixture of oxide and sulfide ores.Thus, in embodiments, the ore may additionally contain laterite ore orconcentrate e.g. saprolite or limonite.

The base metal sulfide ores preferably contain at least one of nickel,cobalt, copper and zinc, as well as at least one platinum group metal(PGM) and/or gold. The value metal content of the ore may vary widely intype and amount, depending on the source of the ore. In particularlypreferred embodiments, the present invention is directed to the recoveryof nickel from base metal sulfide ores, especially nickeliferoussulfide-based ores and mixtures of such ores with related oxide ores.

The process of the present invention may be operated withoutpre-treatment of the base metal sulfide ore. In particular, the processmay be operated without roasting of the ore. However, it may bebeneficial to subject the ore to a grinding or beneficiation step priorto leaching. In particular embodiments of the invention, the ore to betreated may be in the form of a concentrate, and in further embodimentsthe ore to be treated may have been subjected to smelting or other stepsto form a matte. Such steps are known, and are for example discussed inthe references noted above.

Referring to FIG. 1 and FIG. 2, ore 10 in a form as discussed above isfed to a leaching step 12 in which the ore 10 is contacted and leachedwith a lixiviant comprising at least one chloride, hydrochloric acid andat least one oxidant.

The chloride may comprise alkali metal chlorides, alkaline earth metalchlorides, ferric chloride and mixtures thereof. Preferably, thechloride is selected from the group consisting of alkali metalchlorides, magnesium chloride, calcium chloride and mixtures thereof.Preferred examples of alkali metal chlorides include sodium chloride andpotassium chloride. The preferred chlorides are sodium and magnesiumchloride and the chloride may comprise and more preferably consistsessentially of one or both of these chlorides. Most preferably, thechloride comprises or consists essentially of magnesium chloride.Mixtures of chlorides may be used.

The oxidant may comprise alkali metal peroxides, alkaline earth metalperoxides, alkali metal perchlorates, alkaline earth metal perchlorates,ammonium perchlorate, magnesium perchlorate, alkali metal chlorates,alkaline earth metal chlorates, alkali metal hypochlorites, alkalineearth metal hypochlorite, chlorine, hydrogen peroxide and peroxysulfuricacid, and mixtures thereof. Preferably, the oxidant is selected from thegroup consisting of alkali metal peroxides, alkali metal perchlorates,ammonium perchlorate, magnesium perchlorate, alkali metal chlorates,alkaline earth metal chlorates, alkali metal hypochlorites, chlorine,hydrogen peroxide and peroxysulfuric acid, and mixtures thereof.Preferred examples of alkali metal peroxide are sodium peroxide andpotassium peroxide. Preferred examples of alkali metal perchlorates aresodium perchlorate and potassium perchlorate. A preferred example of analkali metal hypochlorite is sodium hypochlorite. Ammonium perchlorate,magnesium perchlorate and peroxysulfuric acid (Caro's acid, H₂SO₅) mayalso be used. Preferred examples of alkali metal chlorates are sodiumchlorate and potassium chlorate. The preferred oxidants are chlorine,sodium hypochlorite, sodium perchlorate and sodium chlorate and theoxidant may comprise and more preferably consists essentially of one ormore of these oxidants.

The leaching step may be conducted in any manner known in the art. Forexample, the leach may be conducted continuously as a co-current step, acountercurrent step or in another manner, or the leaching step may beconducted as a batch step. The leaching step is preferably carried outat atmospheric (ambient) pressure i.e. it is not necessary to conductthe leaching step under pressure. In particular, in accordance with theinstant invention, a leaching step having good reaction kinetics may beconducted at atmospheric pressure. In prior art processes, elevatedpressures are required to obtain reaction kinetics sufficient rapid toenable a commercial process.

It has surprisingly been determined that the formation of sulfate andthe co-dissolution of PGMs and gold may be reduced, and preferablysubstantially reduced, by appropriate selection of the redox potentialand the pH of a lixiviant containing hydrochloric acid and metalchlorides.

The Eh (redox potential versus SHE (standard hydrogen electrode)) may bemaintained in the range of 250-600 mV, preferably from 250-450 mV andmost preferably from 350-450 mV. If the redox potential is less thanabout 250 mV, then the lixiviant is highly reductive and base metalsulfides in the ore will not be leached at an appreciable rate. If theredox potential is higher than about 600 mV, then the PGMs and gold willco-dissolve at an appreciable rate and sulfides leached from the orewill be converted to sulfates at an appreciable rate. Accordingly, it ispreferred to maintain the redox potential sufficiently high to leachbase metal sulfides from the ore but sufficiently low so as toessentially limit the co-dissolution of PGMs and gold and to convertsulfur that is associated with metal sulfides in the ore and is leachedfrom the ore (i.e., sulfide sulfur) to hydrogen sulfide.

The amount of oxidant relates to the redox potential (Eh) of theleaching solution. A particular ore will have an emf. The amount ofoxidant that is present in the lixiviant may be adjusted to obtain adesired redox potential for the lixiviant.

The pH of the lixiviant solution, as measured by conventional equipment,at the end of the leaching operation may be in less than 2.5, althoughit is preferable for the pH to be less than 1.5, more preferably lessthan 1.0 and most preferably in the range 0-0.8. It is to be understoodthat the pH in the leach solution will vary, and might be in the rangeof 0.5-4.0 initially. However, in order to reduce the residence time ofthe leaching step, the pH is preferably maintained in the selected rangefor most (e.g., more than about 50%) of the duration of the leachingstep. At a pH higher than about 0.6 iron commences to precipitate ashematite and magnetic hydroxide (e.g., spinel). The precipitation ofiron substantially increases when the pH is above about 1-1.5.Accordingly, it is preferred to maintain a low pH, especially if thereare significant amounts of PGMs and/or gold in the ore. In one preferredembodiment, if the ore contains amounts of PGMs and/or gold that are notsufficient to warrant a separate recovery step, then it is preferred toconduct the leach so as to leach the iron and to precipitate the iron.The leaching and the precipitation may be conducted in a single step(e.g., reactor). Alternately, the leach may be conducted, the solidsremoved and the leachate then treated to precipitate the iron, therebyproducing a separate iron residue for recovery of iron or disposal.Alternately, if it is desired to recover the PGMs and gold, then it ispreferred to maintain the leached iron in the leachate so that the solidresidue is relatively free of iron, thereby simplifying the recovery ofthe PGMs and gold.

The pH of the lixiviant is reduced by providing a sufficientconcentration of chloride in the lixiviant. Accordingly, the chlorideconcentration of the lixiviant from all sources is adjusted to obtainthe selected pH. The chloride concentration may be in the range 200-500grams of chloride ions per litre of lixiviant solution, preferably200-400 g/L and, more preferably 300-400 g/L. The upper limit on thechloride concentration may depend on the ions present in the leachsolution, especially as a result of leaching of the ore, and resultantformation of complexes. In particular, the chloride concentration ispreferably selected to minimize formation of anionic chloro complexes,especially of ferric iron, FeCl₄ ⁻.

In preferred embodiments of the invention, the chloride ions are derivedfrom metal chlorides and hydrochloric acid, and the chlorideconcentration of, e.g., 200-400g/L, is calculated on the basis of theamount of chloride ions in solutions from both the metal chlorides andthe hydrochloric acid in the lixiviant solution. In particularlypreferred embodiments, the amount of hydrochloric acid may be in therange of 30-150 g/L and the amount of metal chloride (e.g., magnesiumchloride) may be in the range of 80-350 g/L.

The metal chloride/HCl (metal to hydrochloric acid) ratio expressed interms of mass percentage (m/m) in the leach is preferably adjusted tooptimize the leach, based on, for example, the particular ore beingleached and the temperature of the leaching step. The metal/HCl ratio ofthe chloride lixiviant solution may be in the range of 0.1-2.0:1 and,preferably 0.4-1.0:1.

The leach is preferably carried out at a temperature in the range of 75°C. up to the boiling point of the leach solution at ambient pressure,which is about 115° C.

The leach may be carried out with a lixiviant having a low concentrationof hydrochloric acid. Preferably, the hydrochloric acid is added in anamount sufficient to leach all of the base metals and, if desired, theiron and to obtain the selected pH. Therefore, the amount ofhydrochloric acid that is added is preferably about the stoichiometricamount of acid required to leach the selected value metals and maintainthe lixiviant in a selected pH range and, more preferably, a slightexcess (e.g., 105%). Therefore, the amount of acid that is added to thelixiviant may be determined by monitoring the pH of the lixiviant duringthe leaching step and adding additional acid as the pH of the lixiviantincreases above a selected value. The amount of acid that is requiredwill vary depending upon the concentration of value metals in the orethe composition of the ore. In particular, higher amounts of acid willgenerally be required if the ore is more concentrated. Similarly,different sulfides require a different amount of acid during theleaching process. For examples, the overall reactions that can occurduring the leaching are as follows.

-   -   Chalcocite        Cu₂S+2HCl+Cl₂→H₂S+2CuCl₂2Cu₂S+8HCl+O₂→2H₂S+4CuCl₂+2H₂O    -   Covellite CuS+2HCl→H₂S+CuCl₂    -   Bornite 2Cu₅FeS₄+16HCl+5Cl₂→8H₂S+10CuCl₂+2FeCl₃    -   Chalcopyrite        2CuFeS₂+8HCl+Cl₂+4H₂S+2CuCl₂+2FeCl₃4CuFeS₂+20HCl+O₂→8H₂S+4CuCl₂+4FeCl₃+2H₂O    -   Enargite 2Cu₃AsS₄+6HCl+8H₂O→8H₂S+6CuCl₂30 2H₃AsO₄    -   Pentlandite        Ni_(g)S₈+16HCl+Cl₂→8H₂S+9NiCl₂2Ni_(g)S₈+36HCl+O₂→16H₂S+18NiCl₂+2H₂O    -   Subsulfide        Ni₃S₂+4HCl+Cl₂→2H₂S+3NiCl₂2Ni₃S₂+12HCl+O₂→4H₂S+6NiCl₂+2H₂O    -   Spalerite ZnS+2HCl+H₂S+ZnCl₂    -   Cobaltite 4CoAsS+8HCl+6H₂O+5O₂→4H₂S+4CoCl₂+4H₃AsO₄    -   Arsenopyrite FeAsS+3HCl+H₂O+O₂+H₂S+FeCl₃+H₃AsO₄    -   Galena PbS+2HCl+H₂S+PbCl₂

For example, a 30% Ni concentrate will require much more acid (e.g.,10-20 times) than a 3% Cu ore. Accordingly, the concentration ofhydrochloric acid in the lixiviant may be 1-4N and may be less than 18%(mass ratio). Use of such a low concentration of hydrochloric acid, andcontrol of the redox potential Eh and pH, are believed to be importantaspects of the control of the form of the sulfur that is obtained fromthe sulfide in the ore i.e. conversion of the sulfide sulfur to hydrogensulfide, rather than sulfate ion. The amount and type of oxidant usedare factors in the control of Eh.

In particularly preferred embodiments, the lixiviant and leachingconditions are chosen so that base metals are leached from the basemetal sulfide ore but platinum group metals (PGMs) and gold areessentially not leached i.e. the PGMs and gold remain as part of thesolids in the leach and are separated as solids by liquid/solidsseparation, as discussed herein. Control of the leach so that PGMs andgold are separated as solids simplifies subsequent steps for recovery ofvalue metals. It will be appreciated that in the preferred embodiment ofthe invention, the leaching step is controlled so that the sulfidesulfur in the sulfide ore material is converted to hydrogen sulfide,rather than sulfate and that the PGMs and gold are essentially notleached (e.g., less than 10%, preferably less than 5% and morepreferably, less than 1%). However, it will be appreciated that,depending upon the subsequent recovery steps, some PGMs and gold may beleached during the base metal leach step and/or some sulfate may beproduced.

As discussed herein, in the preferred embodiment of the invention theleaching step is controlled so that sulfur in the sulfide ore materialis converted to hydrogen sulfide, rather than sulfate. In thisembodiment, the hydrogen sulfide is stripped from the leach solution,most preferably stripped from the leach solution in a continuous mannerso that the concentration in the leach solution of hydrogen sulfide islow. In preferred embodiments, a gas e.g. air or nitrogen, is fed to theleach solution to aid in the stripping of hydrogen sulfide.

It will be appreciated that at least some of the spent lixiviant ispreferably regenerated and fed to leaching step 12. As shown in FIGS. 1and 2, a recycled chloride lixiviant stream 14, as well as an oxidantstream 16 and a make up stream of chloride 18 are combined to producethe lixiviant that is used during the leaching step. It will beappreciated that leaching step 12 may be conducted in a single reactoror a plurality or reactors in series or parallel. Preferably, leachingstep 12 comprises a single leaching reactor. It will also be appreciatedthat some or all of streams 14, 16 and 18 may be combined in anyparticular order prior to being introduced into the reactor or reactorsin which leaching step 12 is conducted.

The hydrogen sulfide stripped from the leach solution may be treated ina variety of ways, preferably for recovery of elemental sulfur, as willbe apparent to persons skilled in the art. For instance, as shown inFIG. 2, the hydrogen sulfide stream 20 may be subjected to a Clausreaction in step 22. In a typical Claus reaction, an oxygen stream 24 isadded and part of a stream of hydrogen sulfide is oxidized to formsulfur dioxide, and the sulfur dioxide is then reacted with remaininghydrogen sulfide to form elemental sulfur 26. The chemical reaction maybe described as follows:2H₂S+3O₂=>SO₂+H₂O3H₂S+{fraction (3/2)}O₂=>3/n S_(n)+H₂O

The reaction may be carried out in more than one stage, using more thanone catalyst, and high efficiencies of recovery of elemental sulfur e.g.94-97%, may be achieved. The Claus reaction is an exothermic reaction,and energy generated 28 (e.g. in the form of a heated liquid that iscirculated in a heat exchanger) may be sent to the leaching process(e.g. an indirect heat exchanger) or used elsewhere in the process.

In other embodiments, the hydrogen sulfide may be contacted with asolution of a metal that will form a sulfide, especially a solution of acopper salt. The copper salt may be obtained from the spent lixiviantsuch as by selective solvent extraction and stripping. For example asshown by dotted line in FIG. 2, stream 21 of hydrogen sulfide may be maybe combined with the pregnant strip solution produced in extraction step44. Examples of copper salts include cuprous chloride and cupricsulfate. If the recovery process is a so-called stand-alone process,i.e. streams of liquids in the sulfur recovery process are not recycledto the process for leaching and recovery of metal values from thesulfide ore material, then a variety of copper salts could be used.However, if streams from the sulfur recovery process are or might berecycled to the process for leaching and recovery or metal values, thenit is particularly preferred that the copper salt be cuprous chloride.The leaching step is a chloride process, and use of cuprous chloridereduces or eliminates contamination of the leaching step with anionsother than chloride. Contacting hydrogen sulfide with cuprous chloridesolution results in the formation of copper (cuprous) sulfide, which maybe separated in a liquid/solids separation step. The liquid may berecycled and recontacted with hydrogen sulfide. Copper sulfide may beconverted to copper sulfate and elemental sulfur.

The leaching conditions, and especially the lixiviant, redox potentialEh and pH, may be controlled so that at least 50%, preferably at least90% and most preferably at least99% by weight of the sulfide sulfur thatis leached from the sulfide ore material is converted into hydrogensulfide. Under appropriate conditions, at least 99.9% and particularlyat least 99.95%, by weight of the sulfide sulfur in the sulfide orematerial may be converted into hydrogen sulfide. Is

Preferably, the total amount of sulfate formed in the leaching step isless than 1%, more preferably less than 0.1% and most preferably lessthan 0.05%, by weight of the amount of sulfur in the sulfide orematerial 10 that is leached from the ore during leaching step 12. Theformation of hydrogen sulfide, which is stripped, and the low levels ofsulfate, simplifies subsequent steps in the recovery of value metalsand/or recovery and recycling of components of the leach solution.

The leach mixture 30 comprises a value metal-rich solution (leachate) 32and a residue (solids) is in the form of a suspension 34. The leachmixture 30 is fed to a solid/liquid separation step 36 to effectseparation of the leachate 32 from the solids 34. The solids 34 mayinclude unleached ore (e.g., the PGMs and gold) and iron solids,although it would be preferable to maintain iron in solution ifsignificant values of PGMs and/or gold are in the ore feed. Techniquesfor such separation are known e.g. using a pressure or vacuum filter,counter-current decantation or centrifuge.

Solids 34 may comprise a magnetic portion, which may be separated andwhich might be useful for production of ferro-nickel or low-alloystainless steels.

The PGMs and gold in the solids 34 may be recovered by any means knownin the art. Preferably, solids 34 are leached to dissolve the PGMs andgold. Preferably, the lixiviant may be any of those taught herein forleaching the base metals, except that the redox potential is preferablygreater than 700 mV and, more preferably, greater than 800 mV. In aparticularly preferred embodiment, the lixiviant has the samecomposition as that used to leach the base metals from the sulfide oreexcept that the composition has been adjusted to increase the redoxpotential.

As discussed above, it is preferred that the leach be carried out sothat the platinum group metals, and gold and optionally at least aportion of the silver, are not leached so that they may be separatedwith the leach residue, and separated therefrom using known techniques.If some of the platinum group metals and gold are in leachate stream 32that is separated in liquid/solids separation step 36, then some or allof leachate 32 may be subjected to PGM separation step 38 to recover thePGM and gold, and silver if any is present. As shown in FIG. 1, a bleedstream 40 may be removed from leachate 36 and subjected to PGMseparation step 38. A PGM- and gold-poor leachate stream 42 is returnedto leachate stream 32. PGM separation step 38 may be any process knownin the art to remove dissolved PGMs and gold, and optionally silver ifpresent. Preferably, PGM separation step 38 is a cementation step e.g.using copper, zinc or inorganic or organic reductants e.g. sodiumborohydride or hydrazine.

The solids separated in liquid/solid separation step 36 may containcopper sulfide, depending on the Eh of the leach solution, and may berecovered by any process known in the art.

In embodiments of the invention, value metals e.g. nickel, copper, zincand/or cobalt and PGMs and gold may be recovered from the leachate 32 bystandard or other known methods e.g. ion exchange, solvent extraction,electrowinning or sulfide precipitation. Examples are given in FIG. 1and FIG. 2.

In one embodiment for the separation of value metals from the leachate32, copper ions are present in leachate 32 from leaching of coppervalues from ore 12 or from the addition of copper salts e.g. copperchloride, or copper sulfide (Cu₂S) may be formed in the solution e.g.from hydrogen sulfide generated in the leach. As shown in FIGS. 1 and 2,copper could be recovered from the leachate 32 by subjecting leachate 32to a solvent extraction step 44 to obtain a copper reduced leachate 50.lAccordingly, leachate 32 could be contacted with an appropriateextraction solution 46 to obtain a copper rich solution 48 that istreated to recover copper. The extraction solution 48 may then beregenerated and recycled as is known in the art to obtain extractionsolution 46. It is however preferred to form hydrogen sulfide in theleach and strip the hydrogen sulfide from the leach solution, asdiscussed herein.

The copper reduced leachate 50 may also contain iron as Iron chloride.As shown in FIG. 1, the iron may be recovered from leachate 50 by theaddition of magnesium oxide 54 to precipitate an iron oxide (such ashematite or spinel) or a hydrated iron oxide in precipitation step 52.The former are preferable since they are easier to effect solid/liquidseparation. The leach mixture 56 comprises a value metal-rich solution(leachate) 58 and a residue (solids) is in the form of a suspension 60.The leach mixture 56 may be fed to solid/liquid separation step 62 as isknown in the art to effect separation of the leachate 58 from the solids60. Solids 60 may be treated to recover value metals therefrom and/ordisposed as spent solids. The iron may alternately be recovered by beingpyrohydrolysed to form an oxide.

As shown in FIG. 1, a portion of leachate 58 may be removed via stream64 and then subjected to precipitation step 66. For example, magnesiumoxide stream 68 may be combined with stream 64 to form a mixed hydroxideprecipitate (e.g., nickel/cobalt hydroxide) to produce product 70 and avalue metal depleted leachate 72. The remainder of leachate 58 maycontain nickel and cobalt that may be individually recovered in separaterecovery steps. For example, a portion of leachate 58 may be subjectedto a solvent extraction step 74 to obtain a cobalt rich extractionsolution 76 and a cobalt reduced leachate 78. The cobalt rich extractionsolution may be treated to obtain a cobalt containing solution 80 and avalue metal depleted leachate 82. Cobalt reduced leachate 78 may betreated to recover nickel and a value metal depleted leachate stream.

The value metal depleted leachates may be combined and treated toregenerate the lixiviant. In particular, the metal chloride andhydrochloric acid may be regenerated. Further, magnesium oxide used inthe value metal recovery steps may be obtained. Referring to FIG. 1, ableed stream of value metal depleted leachates may be treated in step 84to remove impurities therefrom. A portion of the purified value metaldepleted leachate 86 may then be treated in step 88, such as byhydroxide or sulfide precipitation to obtain metal chloride (e.g.,magnesium chloride) for recycle. A portion of the purified value metaldepleted leachate 86 may be subjected to pyrohydrolysis step 90 toobtain magnesium oxide stream 54, 68

The magnesium oxide reduced leachate 92 produced by hydrolysis step 90comprises HCl that may be subjected to additional evaporation steps 94to obtain recycle HCl. Off-gases from pyrohydrolysis may be used inpre-evaporation (not shown), to enrich the solution in HCl and reduceenergy costs. However, the degree of partial or pre-evaporation may bereduced, or even eliminated, by feeding gaseous hydrogen chloride to thesolution. The hydrogen chloride may be formed from chlorine. In thismanner, energy required for evaporation of water may be reduced oreliminated.

In the alternate embodiment shown in FIG. 2, leachate 50 is treated intwo or three purification steps 100 to recover value metals that areprecipitated from solution by the addition of magnesium oxide, which isprovided by stream 102. After each purification step 100, a treatedleachate 104 is subjected to solid/liquid separation step 106 to obtaina solid 108, which may be sent for disposal or for further processing toisolate the value metal, and a value metal reduced leachate 110.Subsequently, the treated leachate is subjected to a nickel/cobaltrecovery step 112 to obtain a value metal reduced leachate 114, whichmay be recycled by pyrohydrolysis step 90, and a mixed nickel/cobalthydroxide product, which may then be subjected to further processing toisolate the value metals. Alternately, the leachate may be treatedsequentially to produce a nickel-containing product andcobalt-containing product.

It will be appreciated that by sequentially adding additional amounts ofmagnesium oxide, the pH of the leachate may be sequentially increased soas to precipitate a particular metal or group of metals. It will also beappreciated that the pH of the leachate may be adjusted by variousmeans, including the addition of different pH adjustment agents (e.g.bases). An advantage of the use of magnesium oxide is that the requiredamount of magnesium oxide may be produced by the process and theaddition of magnesium oxide does not add any additional ions in theleachate, which may require the use of additional treatment steps.

The lixiviant, especially redox potential (Eh), is controlled to effectconversion of sulfide in the sulfide ore material fed to the leachingstep into hydrogen sulfide, rather than sulfate ion. The hydrogensulfide is preferably stripped from the leaching step as gaseoushydrogen sulfide. In the embodiment in which hydrogen sulfide is formed,formation of sulfate may be reduced to very low levels e.g. 0.05% byweight or lower based on the total amount of sulfide sulfur leached fromthe ore, as exemplified herein, which facilitates separation of valuemetals in subsequent steps in the process.

Thus, in preferred embodiments of the present invention e.g. as shown inFIG. 1 and FIG. 2, the present invention provides for the use ofmixtures of magnesium chloride, at least one oxidant and hydrochloricacid in the leach step. Sulfide in the ore may be separated as elementalsulfur, or most preferably as hydrogen sulfide. The dissolution of ironmay be controlled and minimized, without requiring expensivepre-treatment or post-treatment steps by adjustment of chlorideconcentration, pH, kinetics, redox and/or temperature. For example,lower leaching temperatures, lower chloride concentrations and a higherpH decrease the tendency of iron to be leached. Therefore, thetemperature, chloride concentration and pH conditions for the leach maybe selected, in part, based on the amount of iron to be leached.Subsequent to the iron being leached, the pH may be increased,preferably above 1.5 to precipitate the leach iron. The leach residuemay be maintained in a form that is readily filterable. As discussedherein, the process is preferably controlled so that hydrogen sulfide isformed during leaching, and stripped from the leach solution prior tosubsequent liquid/solids separation of the leach solution.

In the process of the present invention, the metal chloride/HCl ratioe.g. metal/HCl ratio and the amount and type of oxidant in the leachstep may be adjusted to reflect any specific requirements orcharacteristics of the process and ore fed to the process. In someinstances, all of the chloride ion in the leach solution may be suppliedfrom, for example, recycled magnesium chloride.

The leaching of the base metals may be conducted continuously in atleast one stirred tank reactor. Alternately, at least two reactors maybe used, the first for addition of base metal sulfide ore and the secondfor removal of the leached iron (e.g., increasing the pH to precipitatethe iron either as part of the solid residue from the leach or, as shownin FIG. 1, in a separate precipitation step 52 downstream fromsolid/liquid separation step 36). Should there be significant PGM and/orgold values in the feed, it will be preferable to maintain the iron insolution until after separation of the leach residue, which will be aPGM and/or gold concentrate. If there are no PGM and/or gold valuespresent, then the iron will be precipitated into the leach residue byadjustment of the parameters indicated above. Three or more reactors maybe more optimal. Process control may be effected by the rates ofaddition of base metal sulfide ore and/or lixiviant solution to theprocess, but it may be preferable to control the process using pH andredox potential Eh. As discussed above, the leaching may also beconducted batch, co-current or countercurrent, in whole or in part. Itwill also be appreciated that any of the downstream process steps may beconducted on a continuous or a batch basis.

An increase has been recognized in the activity of HCl when salts suchas NaCl, CaCl₂ and MgCl₂ are added to dilute solutions of HCl. Withoutbeing limited by theory, the increase in the reactivity of HCl isunderstood to be a function of chloride ion concentration, especially ofmagnesium chloride. Magnesium chloride has a high hydration rate, whichis believed to cause substantially increased activity of hydrogen ionsin the lixiviant solution. However, as illustrated by comparativeexperiments below, especially Example II, use of constant chlorideconcentrations in hydrochloric acid leaching of nickeliferous ore canlead to widely different levels of extraction of nickel.

00] The process provides for removal of sulfide sulfur derived from thesulfide ore as hydrogen sulfide, rather than the formation of sulfates.The preferred embodiment of removal of sulfur as hydrogen sulfidesimplifies and/or allows for alternate steps for separation and recoveryof value metals subsequent to the leaching step, because sulfate ispresent in not more than very minor amounts, as exemplified herein. Inaddition, leaching conditions, especially pH, redox potential (Eh) andchloride concentration, may be controlled thereby providing for controlof leaching of value metals, formation of chloride complexes andextraction of iron, in addition to control so that sulfide sulfur isconverted to hydrogen sulfide. The process of the present invention doesnot require pre-treatment of the base metal sulfide ore prior to theleaching step.

A particular advantage of the process of the present invention is thatboth high rates of extraction of value base metals and removal of sulfurhydrogen sulfide may be obtained in a leaching step that operates atatmospheric pressure. The use of low concentrations of hydrochloric acidand the use of high levels of chloride in the lixiviant, preferablymagnesium chloride, at the selected redox potential results in theformation of hydrogen sulfide. The hydrogen sulfide is stripped from theleach solution, and results in very low amounts of sulfate in leachateand solids from the leach solution. This has significant economicadvantages in subsequent steps for recovery of value metals and PGMs andgold. In addition, the use of the low concentrations of hydrochloricacid does not effect leaching of PGMs and gold from the ore, which alsosimplifies subsequent steps in the recovery of value metals from theleachate. The use of atmospheric pressure results in substantialeconomic advantages, especially in capital costs. The use of chloridechemistry offers advantages in operating and capital costs of theprocess. The leaching agent is regenerated and recycled, preferablyusing a pyrohydrolysis step with additional hydrochloric acid beingformed from chlorine if required. Magnesium chloride is the preferredchloride, as it is more readily recycled to the leaching step.Additionally, the use of magnesium chloride with hydrochloric acid andoxidant as lixiviant is preferred.

While not being bound by any theory, the high activity of H⁺ ions in thehigh strength chloride solutions, especially magnesium chloridesolutions, is believed to enable use of lower concentrations ofhydrochloric acid to effect leaching of value metals, and in embodimentsit is believed that the amount of acid required may be only marginallyhigher than the stoichiometric amount of acid. The high activity of theproton, H⁺, in high concentration chloride solutions permits even smallamounts of acid to act as though it were highly concentrated, andtherefore has a driving force and hence very little excess overstoichiometric acid is required. Where the proton activity is not sohigh, then considerable excess acid is required to drive the leachingreaction. There is lower water activity in the chloride solutions, whichis believed to result in lower concentrations of iron in solution, butwith the relatively low amounts of iron in the ores percentageextraction is still high. The presence of magnesium ions in magnesiumchloride solutions is believed to reduce dissolution of magnesium as aresult of common ion effects. Use of magnesium chloride permits recycleof both hydrochloric acid and caustic (highly reactive) magnesia, bothof which may be used in the process. The high proton activity achievableat the low acid concentrations herein permits the process to be operatedunder conditions that cause formation of hydrogen sulfide rather sulfateion, so-called reductive leaching conditions at low redox potential. Itis also believed that the low redox potential used in the process notonly results in the formation of hydrogen sulfide instead of sulfate butalso is not conducive to leaching of PGMs and gold. Both of theseaspects are advantages of the process of the present invention.

The present invention is illustrated by the following examples.

EXAMPLE I

A series of comparative laboratory-scale leaching experiments werecarried out using a base metal sulfide ore concentrate that was -amixture of pyrrhotite, pentlandite and chalcopyrite. The ore concentratehad the following analysis: Ni (18.65%), Cu (1.38%), Co (0.19%) and Fe(26.6%). The leach solution (approximately 500 mL) was a hydrochloricacid (2N in Runs 1 and 3, 4N in Run 2) solution containing 20 w/w ofsolids. Ferric chloride (FeCl₃.6H₂O ) was added to each leach solution,so that the total chloride ion content (from HCl and ferric chloride)was 230 g/L. The temperature of the solution was 95° C. and the leachingtime was 4 hours. The redox potential (Eh) was measured in mV. The pHwas less than 0.

Oxidant was not added in Runs 1 and 2. In Run 3, chlorine gas wasbubbled through the leach solution at a rate of 0.5 mL/min.

The leached solution was subjected to a liquid/solids separation step.The washed solids obtained were subjected to analysis for the content ofnickel, iron, cobalt and copper, and the liquid was subjected toanalysis for nickel. The extraction of each metal was then calculated.

The results obtained, expressed as percentages based on the concentratefed to the leach solution, are shown in Table I. Table I shows the redoxpotential, in mV, at the end of the leach. TABLE I Run 1 Run 2 Run 3Redox potential (Eh) 170 235 510 Ni extraction (solids) 36 23 84 Niextraction (liquid) 47 23 65 Fe (solids) 70 46 68 Co extraction (liquid)70 16 56 Cu extraction (liquid) 75 15 63

The results show that leaching of the sample using leach solution ofhydrochloric acid and ferric chloride resulted in poor leaching ofnickel (below 50% extraction) except when chlorine was bubbled throughthe leach solution (Run 3). The leach solution of Run 3 exhibited ahigher redox potential.

EXAMPLE II

In further comparative experiment, the procedure of Example 1 wasrepeated using 2N hydrochloric acid, except that for Run 4, the metalchloride was magnesium chloride (MgCl₂.6H₂O). The total chloride ioncontent was 300 g/L. The pH was <0.

The results obtained are given in Table II. TABLE II Run 4 Redoxpotential (Eh) 410 Ni extraction (solids) 64 Ni extraction (liquid) 66Fe (solids) 74 Co extraction (liquid) 58 Cu extraction (liquid) 79

The results show that the use of magnesium chloride (Run 4) improved theextraction of nickel from 47% to 66% as against Run 1, and wasequivalent to ferric chloride and chlorine (Run 3) in the extraction ofiron, cobalt and copper and slightly poorer in the extraction of nickel

EXAMPLE III

The procedure of Example II was repeated, except that an oxidant wasadded to the leach solution, in order to illustrate aspects of thepresent invention. Thus, in each Run, the leach contained both magnesiumchloride and an oxidant. The amounts, based on 500 mL of solution, areshown in Table III; chlorine was fed into the leach solution in theamount shown. In Run 5, the hydrochloric acid was 2N and the amount oftotal chloride ion concentration was 300 g/L. In Runs 6-9, thehydrochloric acid was 4N and the amount of total chloride ionconcentration was 400 g/L. The pH was <0.

The results obtained are shown in Table III. TABLE III Run 5 Run 6 Run 7Run 8 Run9 Oxidant NaClO₃ NaClO₃ NaClO₃ NaClO₃ Cl₂ Amount of oxidant 117g 0.25 g  0.5 g 0.75 g  0.5 mL/min MgCl₂.6H₂O 112 g  370 g  370 g  370 g 370 g Redox pot. (Eh) 340 430 440 420 510 Ni extrn. (solids) 54 95 9692  98 Ni extrn. (liquid) 57 78 86 88  92 Fe (solids) 32 91 92 88  98 Coextrn. (liquid) 55 90 88 85  96 Cu extrn. (liquid) 64 48 28 24  99

The results show that high extractions of nickel, especially in excessof 90%, may be obtained using leach solutions containing hydrochloricacid, magnesium chloride and an oxidant. In addition, while theextraction of nickel increases at redox potentials above 250 mV, theextent of the extraction of nickel increases substantially at redoxpotentials above 350 mV.

EXAMPLE IV

A series of laboratory-scale leaching experiments were carried out usinga base metal sulfide ore concentrate that was a mixture of pyrrhotite,pentlandite and chalcopyrite. The feed ore was received at 100% -100mesh feed size. The ore concentrate had the following analysis: Ni(18.90%), Cu (1.52%), Co (0.28%) and Fe (29.0%). The leach solution (1L)was a 4N hydrochloric acid solution containing 5% w/v of solids. Theleach solution had a total chloride ion concentration of 400 g/L,obtained from hydrochloric acid and magnesium chloride (258.2 g/L) asCl—. The leaching temperature was 95° C. and the leaching time was 4hours. The redox potential (Eh) was measured in mV. Chlorine was bubbledthrough the leach solution as oxidant. The pH was <0.

Cl₂ was bubbled through the leach solution to strip hydrogen sulfideformed during the leach from the leach solution.

The leached solution was subjected to a liquid/solids separation step.The solids (washed) and liquid obtained were subjected to analysis forthe content of nickel, iron, copper and cobalt. The extraction of eachmetal was then calculated. The liquid was also subjected to analysis forsulfate.

The results obtained, expressed as percentages based on the analysis ofthe solids and liquid for each metal, are shown in Table IV. Theanalysis for sulfate is reported as both g/L of sulfate and percentageremoval of sulfide sulfur as hydrogen sulfide. Table IV shows the redoxpotential, in mV, at the end of the leach. In each run reported in TableIV, visual examination of the solids showed no evidence of elementalsulfur. TABLE IV Run 10 Run 11 Run 12 Redox potential (Eh) 190 186 295Ni extraction (%) 87.9 59.4 67.6 Fe extraction (%) 83.5 72.3 73.2 Cuextraction (%) 34.8 50.0 61.7 Co extraction (%) 72.0 38.8 50.1 Sulfate(liquid, g/L) 0.011 0.020 0.010 Removal of sulfide sulfur (%) >99 >99>99

The results show that in each of Runs 10-12 at least 99% of the sulfur,which was in the form of sulfide in the sulfide ore concentrate fed tothe leach, had been removed during the leaching step. The amount ofsulfate ion in the liquid obtained was very low.

EXAMPLE V

The procedure of Example IV was repeated, except that chlorine was notbubbled through the leach solution. The oxidant used was sodiumperchlorate, which was added in amounts as follows: Runs 13 and 14-20kg/tonne of concentrate sample; Run 15-10 kg/tonne of concentratesample. In addition to the sodium perchlorate, in Runs 14-15 oxygen wasbubbled through the leach solution (1L) at a rate of 100 mL/min. Sodiumperchlorate was the only oxidant used in Run 13. The pH was <0.

The results obtained are shown in Table V. In each run reported in TableV, visual examination of the solids showed no evidence of elementalsulfur. TABLE V Run 13 Run 14 Run 15 Redox potential (Eh) 405 250 365 Niextraction (%) 96.1 51.2 97.3 Fe extraction (%) 90.4 65.5 92.2 Cuextraction (%) 81.7 52.4 79.3 Co extraction (%) 83.2 33.4 83.5 Sulfate(liquid, g/L) 0.004 0.042 0.007 Removal of sulfide sulfur (%) >99 >99>99

The results for removal of sulfide sulfur using sodium perchlorate orsodium perchlorate/oxygen as oxidant are similar to those of Example IVin which chlorine was used as oxidant. High extractions of nickel, inexcess of 90%, were obtained in Runs 13 and 15. This example shows thatsodium perchlorate may be used as oxidant instead of chlorine. At least99% of the sulfide sulfur was removed in all Runs.

EXAMPLE VI

The procedure of Example V was repeated, except that the only oxidantused was sodium hypochlorite. A solution of sodium hypochlorite (5.25%w/v) was added as follows: Run 16—560 L/tonne of concentrate sample; Run17—160 L/tonne of concentrate sample; and Run 18—960 L/tonne ofconcentrate sample. The pH was <0.

The results obtained are shown in Table VI. In each run reported inTable VI, visual examination of the solids showed no evidence ofelemental sulfur. TABLE VI Run 16 Run 17 Run 18 Redox potential (Eh) 398295 190 Ni extraction (%) 98.2 98.8 42.2 Fe extraction (%) 91.9 89.556.6 Cu extraction (%) 85.8 75.7 43.2 Co extraction (%) 86.2 83.4 23.1Sulfate (liquid, g/L) 0.023 0.017 0.018 Removal of sulfide sulfur(%) >99 >99 >99

The results for removal of sulfide sulfur using sodium hypochlorite asoxidant are similar to those of Examples IV and V, in which chlorine,sodium perchlorate and sodium perchlorate/oxygen were used as oxidant.High extractions of nickel, in excess of 90%, were obtained in Runs 16and 17. In Run 18, the high concentration of sodium hypochlorite isbelieved to have resulted in re-precipitation of dissolved metals, butthe removal of sulfide sulfur as hydrogen sulfide gas was still high.

EXAMPLE VII

A series of laboratory-scale leaching experiments were carried out usinga material identified as an anode slime material and oxidants such assodium hypochlorite, sodium chlorate and oxygen. The anode slimematerial had the following analysis: Ni (5.1%), Cu (1.1%), Fe (1.2%), Au(7.01 g/t), Ag (2.20g/t), Pt (24.40g/t), Pd (87.90g/t) and Rh (3.50g/t).The sulfur content was 91.9 wt %, with the content of elemental sulfurbeing 80 wt %. The leach solution (1L) was a 4N hydrochloric acidsolution containing 5% w/v of solids. The leach solution had a totalchloride ion concentration of 400 g/L, obtained from hydrochloric acidand magnesium chloride (258.2 g/L). The leaching temperature was 95° C.and the leaching time was 6 hours (Run 19 and Run 20). The redoxpotential (Eh) was measured in mV. The pH was <0.

Air was bubbled through the leach solution to strip hydrogen sulfideformed during the leach from the leach solution.

The leached solution was subjected to a liquid/solids separation step.The solids (washed) and liquid obtained were subjected to analysis forthe content of nickel, iron, copper, gold, silver, platinum, palladiumand rhodium. The extraction of each metal was then calculated.

The results obtained, expressed as percentages based on the analysis ofthe solids, are shown in Table VII, together with the weight loss of theanode slime after leaching. Table VII shows the redox potential, in mV,at the end of the leach. In Run 19 the oxidant was sodium hypochloriteat 120 L of 5.25 w/v % NaOCl per tonne of anode slime sample, and in Run20 the oxidant was sodium chlorate added at 160 kg/tonne of ore. TABLEVII Run 19 Run 20 Redox potential (Eh) 550 >850 Weight loss (%) 18.818.8 Ni extraction (%) 86 107 Fe extraction (%) 117 132 Cu extraction(%) 92 104 Au extraction (%) n/a 64 Ag extraction (%) n/a >82 Ptextraction (%) n/a 56 Pd extraction (%) n/a 62 Rh extraction (%) n/a 64n/a = not analyzed

The results show that value metals can be extracted from a secondaryfeed such as an anode slime material. The results also demonstrate thatat higher redox levels, PGMs and gold may also be leached.

EXAMPLE VIII

A sample of a copper sulfide/oxide ore containing 1.275 wt % copper,2.61 wt % iron and 0.83 wt % sulfur was subjected to a leach using aleach solution of hydrochloric acid (4N) and magnesium chloride, at atotal chloride ion concentration of 400 g/L, that additionally contained20 kg/tonne s of sodium chlorate. The leach was conducted for 4 hours at95-105° C. The pH was <0. The ore was minus 150 mesh and was used at 5wt % solids. The residue from the leach was analyzed for copper, ironand sulfur.

It was found that 97.3% of the copper, 69.4% of the iron and 54.6% ofthe sulfur in the ore had been leached, thereby demonstrating theleaching of a mixed sulfide/oxide ore of copper. The weight loss inleaching was 10.2 % and the terminal redox potential was greater than450 mV.

EXAMPLE IX

The procedure of Example VIII was repeated, using the same ore, exceptthat the leach was conducted on 10 wt % solids and the sodium chloratewas used in an amount of 10 kg/tonne. In addition, the resultant leachsolution, not the residue was analyzed. In this test (Run 21), it wasfound that 96.5% of the copper and 88.8% of the iron had been leachedinto solution, thereby illustrating the invention. The weight loss inleaching was 17.8 % and the terminal redox was 450 mV. The pH <0.

The procedure of Example VIII was repeated using the same ore, exceptthat the oxidant used was 48 kg/tonne of hydrogen peroxide and the leachwas conducted for 6.5 hours. Analysis of the leach solution showed that92.4 wt % of the copper and 88.5 wt % of the iron had been leached intosolution, thereby showing that hydrogen peroxide was less effective thansodium chlorate as oxidant. The weight loss in leaching was 11.8 % andthe terminal redox was greater than 450 mV. The pH was <0.

In a comparative test (Run 22), the procedure of Example VIII wasrepeated using the same ore but without addition of any oxidant. Theleach time was increased from 4 hours to 12 hours. Analysis of the leachsolution showed that 96.2 wt % of the copper and 82.1 wt % of the ironhad been leached into solution, thereby showing that a substantiallyincreased leach time was required in the absence of oxidant. The weightloss in leaching was 11.0 % and the terminal redox was 420 mV. The pHwas <0.

EXAMPLE X

A sample of a polymetallic sulfide feed material containing 0.93 wt %copper, 44.9 wt % iron, 0.66 wt % nickel, 0.06 wt % cobalt, 23.9 wt %sulfur, 0.50 g/t platinum, 1.79 g/t palladium, 0.02 g/t rhodium, 0.03g/t gold and 1.6 g/t silver was subjected to a leach using a leachsolution of hydrochloric acid (4N) and magnesium chloride, at a totalchloride ion concentration of 400 g/L, that additionally contained 20kg/tonne of sodium chlorate. The leach was conducted for 4 hours at100-105“C. The feed material was used at 5 wt % solids. The pregnantleach solution obtained from the leach was analyzed for copper, iron,nickel and cobalt. The pH was <0.

It was found that 97.1 % of the copper, 98.1 % of the iron, 93.2% of thenickel and 71.5% of the cobalt in the polymetallic sulfide feed materialhad been leached into solution, thereby demonstrating the leaching of apolymetallic sulfide material. The weight loss in leaching was 82.8% andthe terminal redox potential was 480 mV. Only 0.1% Pt, 0.02% Pd, 2.6%Rh, 0.6% Au and 16.3% Ag were extracted in this controlled oxidationleach test.

1. A process for leaching a value metal from a sulfide ore materialcontaining said value metal, said sulfide ore material, comprising thestep of leaching the sulfide ore material at atmospheric pressure with alixiviant comprising hydrochloric, a chloride selected from the groupconsisting of alkali metal chlorides, magnesium chloride and calciumchloride, and mixtures thereof, and an oxidant selected from the groupconsisting of alkali metal peroxide, alkali metal perchlorate, ammoniumperchlorate, magnesium perchlorate, alkali metal chlorate, alkalineearth metal perchlorate, chlorine, alkali metal hypochlorite, hydrogenperoxide and peroxysulfuric acid, and mixtures thereof to obtain aleachate and a solid residue.
 2. The process of claim 1 furthercomprising selecting a sulfide ore material that contains at least onevalue metal selected from the group consisting of nickel, copper, zincand cobalt, and mixtures thereof.
 3. The process of claim 2 furthercomprising selecting a sulfide ore material that additionally containsat least one of gold and a platinum group metal.
 4. The process of claim2 further comprising selecting a sulfide ore material that comprises abase metal sulfide ore, a concentrate of a base metal sulfide ore, amatte obtained from a base metal sulfide ore, tailings from theprocessing of a base metal sulfide ore and mixtures thereof.
 5. Theprocess of claim 4 further comprising selecting pyrrhotite, pentlandite,chalcopyrite, pyrite, arsenopyrite and sphalerite, and mixtures thereofas the base metal sulfide.
 6. The process of claim 2 further comprisingselecting sodium chloride, potassium chloride, magnesium chloride andcalcium chloride, and mixtures as the chloride.
 7. The process of claim5 further comprising selecting magnesium chloride as the chloride. 8.The process of claim 2 further comprising selecting chlorine, sodiumchlorate, hydrogen peroxide, sodium hypochlorite and sodium perchlorateand mixtures thereof as the oxidant.
 9. The process of claim 5 furthercomprising selecting sodium chlorate, chlorine and mixtures thereof asthe oxidant.
 10. The process of claim 2 further comprising adjusting thepH and the redox potential during the leach so that the pH is less than2.5 and the redox potential is in the range of 200 to 600 mV.
 11. Theprocess of claim 10 further comprising adjusting the redox potentialduring the leach so that the redox potential of the leach solution is inthe range 250-450 mV.
 12. The process of claim 10 further comprisingselecting the concentration of magnesium chloride to be at least 200g/L.
 13. The process of claim 12 further comprising selecting the totalconcentration of chloride ions to be in the range of 200-500 g/L. 14.The process of claim 13 further comprising selecting the totalconcentration of chloride ions to be in the range of 200-400 g/L, saidtotal concentration being formed essentially from magnesium chloride andhydrochloric acid.
 15. The process of claim 14 further comprisingselecting the amount of hydrochloric acid to be in the range of 30-150g/L.
 16. The process of claim 10 further comprising conducting the leachat a temperature in the range of from 75° C. to the boiling point of thesolution at ambient pressure.
 17. The process of claim 13 furthercomprising adjusting the pH so that, at the end of the leach, the pH isless than 1.5.
 18. The process of claim 14 further comprising adjustingthe pH so that, at the end of the leach, the pH is less than
 1. 19. Theprocess of claim 3 in which the leachate and a solid residue issubjected to a solids/liquid separation step and the solid residue issubjected to a further leaching step to recover at least a portion ofthe gold and at least one of the platinum group metal.
 20. The processof claim 19 in which the solid residue is also treated for separation ofa magnetic fraction.
 21. The process of claim 1 further comprisingselecting the ore or concentrate to comprise values of nickel and iron,and one or more of cobalt and copper.
 22. The process of claim 1 furthercomprising: a) subjecting the leachate to a series of value metalrecovery steps and obtaining a value metal depleted leachate and b)treating the value metal depleted leachate to recycle at least a portionof the hydrochloric acid and chloride.
 23. The process of claim 22wherein step (b) of claim 22 comprises pyrohydrolysis
 24. The process ofclaim 23 wherein step (a) of claim 22 includes at least oneprecipitation step using a base.
 25. The process of claim 24 furthercomprising selecting magnesium oxide as the base.
 26. The process ofclaim 25 wherein step (b) of claim 22 includes producing magnesium oxidefrom the value metal depleted leachate.
 27. A process for leaching avalue metal from a sulfide source material containing said value metalwith a lixiviant comprising hydrochloric, a metal chloride, and anoxidant, the leaching with said lixiviant being controlled so that atleast about 50% of sulfide sulfur that is leached from the sulfidesource material is converted to hydrogen sulfide during said leaching toobtain a leachate and a solid residue.
 28. The process of claim 27 inwhich the leaching is controlled so that at least about 90% of sulfidethat is leached from the sulfide source material is converted tohydrogen sulfide during said leaching.
 29. The process of claim 27 inwhich the leaching is controlled so that at least about 99% of sulfidesulfur that is leached from the sulfide source material is converted tohydrogen sulfide during said leaching.
 30. The process of claim 27further comprising selecting the sulfide source material to contain atleast one value metal selected from the group consisting of nickel,copper, zinc and cobalt, and mixtures thereof and to additionallycontain at least one of gold and platinum group metal and the leachingis controlled so that the gold and platinum group metals are essentiallynot leached.
 31. The process of claim 27 further comprising selectingthe sulfide source material to contain at least one value metal selectedfrom the group consisting of nickel, copper, zinc and cobalt, andmixtures thereof and to additionally contain at least one of gold andplatinum group metals and the leaching is controlled so that less than10 weight percent of the gold and platinum group metals are leached. 32.The process of claim 27 wherein the leaching is controlled by adjustmentof the pH and the redox potential.
 33. The process of claim 27 furthercomprising adjusting the redox potential of the lixiviant to be from 250to 600 mV.
 34. The process of claim 33 further comprising adjusting thepH of the lixiviant at the end of the leach to be less than 2.5.
 35. Theprocess of claim 34 further comprises adjusting the lixiviant so as tohave a concentration of chloride ions of from 200-500g/L.
 36. Theprocess of claim 34 further comprising selecting the amount ofhydrochloric acid that is added to obtain a selected pH.
 37. The processof claim 27 further comprising selecting the sulfide source materialfrom a base metal sulfide ore or a material derived from a base metalsulfide ore.
 38. The process of claim 37 further comprising selectingthe base metal sulfide ore from the group consisting of pyrrhotite,pentlandite, chalcopyrite, pyrite, arsenopyrite and sphalerite, andmixtures thereof.
 39. The process of claim 27 further comprisingselecting the chloride from the group consisting of alkali metalchlorides, magnesium chloride and calcium chloride, and mixturesthereof, and selecting the oxidant from the group consisting of alkalimetal peroxide, alkali metal perchlorate, ammonium perchlorate,magnesium perchlorate, alkali metal chlorate, alkaline earth metalperchlorate, chlorine, alkali metal hypochlorite, hydrogen peroxide andperoxysulfuric acid, and mixtures thereof.
 40. The process of claim 27in which the chloride comprises an alkali metal chloride, magnesiumchloride and calcium chloride, and mixtures thereof, and the oxidantcomprises an alkali metal peroxide, alkali metal perchlorate, ammoniumperchlorate, magnesium perchlorate, alkali metal chlorate, alkalineearth metal perchlorate, chlorine, alkali metal hypochlorite, hydrogenperoxide and peroxysulfuric acid, and mixtures thereof.
 41. The processof claim 27 in which the chloride comprises sodium chloride, potassiumchloride, magnesium chloride and calcium chloride, and mixtures thereof.42. The process of claim 27 in which the chloride essentially comprisesmagnesium chloride.
 43. The process of claim 41 in which the oxidantcomprises chlorine, sodium chlorate, sodium hypochlorite, sodiumperchlorate and mixtures thereof.
 44. The process of claim 43 in whichthe redox potential of the lixiviant is 250-600 mV.
 45. The process ofclaim 44 in which the concentration of magnesium chloride is at least200 g/L.
 46. The process of claim 27 further comprising conducting theleach at atmospheric pressure.
 47. The process of claim 28 furthercomprising: a) subjecting the leachate to a series of value metalrecovery steps and obtaining a value metal depleted leachate and b)treating the value metal depleted leachate to recycle at least a portionof the hydrochloric acid and chloride.
 48. The process of claim 47wherein step (b) of claim 47 comprises pyrohydrolysis
 49. The process ofclaim 48 wherein step (a) of claim 47 includes at least oneprecipitation step using a base.
 50. The process of claim 49 furthercomprising selecting magnesium oxide as the base.
 51. The process ofclaim 50 wherein step (b) of claim 47 includes producing magnesium oxidefrom the value metal depleted leachate.
 52. A process for leaching avalue metal from a sulfide ore comprising leaching the sulfide ore witha lixiviant comprising hydrochloric acid, a metal chloride, and anoxidant to obtain a leachate and a solid residue, wherein the redoxpotential is maintained sufficiently low to essentially not leachplatinum group metals and gold from the sulfide ore material and toessentially convert the sulfide sulfur which is leached from the sulfideore to hydrogen sulfide.
 53. The process of claim 52 in which thesulfide ore comprises a base metal sulfide ore, a concentrate of a basemetal sulfide ore, tailings from the processing of a base metal sulfideore and mixtures thereof
 54. The process of claim 53 wherein the sulfideore is unroasted prior to being leached.
 55. The process of claim 54further comprising conducting the leach at atmospheric pressure.
 56. Theprocess of claim 55 wherein at least 90% by weight of the sulfide sulfurthat is leached from the sulfide ore material is converted to hydrogensulfide.
 57. The process of claim 56 further comprising the step oftreating at least some of the hydrogen sulfide to obtain elementalsulfur.
 58. The process of claim 56 in which the pH of the lixiviant atthe end of the leach is less than 2.5
 59. The process of claim 52 inwhich the pH of the lixiviant at the end of the leach is less than 1.560. The process of claim 52 in which the pH of the lixiviant at the endof the leach is less than
 1. 61. The process of claim 52 in which the pHof the lixiviant at the end of the leach is less than
 0. 62. The processof claim 58 in which the redox potential (Eh) is in the range of 250 to600 mV.
 63. The process of claim 59 in which the redox potential (Eh) isin the range of 250 to 450 mV.
 64. The process of claim 52 in which theredox potential (Eh) is in the range of 250 to 350 mV.
 65. The processof claim 62 in which the lixiviant has a concentration of chloride ionsof at least 200 g/L.
 66. The process of claim 52 in which the lixivianthas a concentration of chloride ions is in the range of 200-500 g/L. 67.The process of claim 52 in which the lixiviant has a concentration ofchloride ions of 300-400 g/L
 68. The process of claim 65 in which theconcentration comprises the chloride ions obtained from the chloride andthe chloride ions obtained from the hydrochloric acid.
 69. The processof claim 66 in which the amount of hydrochloric acid is in the range of30-150 g/L.
 70. The process of claim 52 in which the chloride comprisesone or more of alkali metal chlorides, magnesium chloride and calciumchloride, and mixtures thereof, and the oxidant comprises one or more ofalkali metal peroxide, alkali metal perchlorate, ammonium perchlorate,magnesium perchlorate, alkali metal chlorate, alkaline earth metalperchlorate, chlorine, alkali metal hypochlorite, hydrogen peroxide andperoxysulfuric acid, and mixtures thereof.
 71. The process of claim 52in which chloride is selected from the group consisting of alkali metalchlorides, magnesium chloride and calcium chloride, and mixturesthereof, and the oxidant is selected from the group consisting of alkalimetal peroxide, alkali metal perchlorate, ammonium perchlorate,magnesium perchlorate, alkali metal chlorate, alkaline earth metalperchlorate, chlorine, alkali metal hypochlorite, hydrogen peroxide andperoxysulfuric acid, and mixtures thereof.
 72. The process of claim 52in which the leachate and a solid residue is subjected to asolids/liquid separation step and the solid residue is subjected to afurther leaching step to recover at least one of the platinum groupmetals and gold.
 73. The process of claim 52 further comprising: a)subjecting the leachate to a series of value metal recovery steps andobtaining a value metal depleted leachate and b) treating the valuemetal depleted leachate to recycle at least a portion of thehydrochloric acid and chloride.
 74. The process of claim 73 wherein step(b) of claim 73 comprises pyrohydrolysis
 75. The process of claim 74wherein step (a) of claim 73 includes at least one precipitation stepusing a base.
 76. The process of claim 75 further comprising selectingmagnesium oxide as the base.
 77. The process of claim 76 wherein step(b) of claim 73 includes producing magnesium oxide from the value metaldepleted leachate.
 78. The process of claim 73 further comprisingtreating the leachate to precipitate iron.
 79. The process of claim 78wherein the leachate is treated by increasing the pH of the leachate,subsequent to removal of the residual solids to, to precipitate iron.